The present invention relates to a method of treating zinc concentrates for recovering zinc, and especially to a method for recovering zinc which includes successive smelting (and oxidizing) and fuming steps. During the smelting and oxidizing stage, zinc concenrate is oxidized and smelted by supplying gas containing air or oxygen into it, whereby sulphur dioxide is exhausted and the zinc oxide thus produced is slagged. The slag produced in the smelting and oxidizing stage is then reduced in a first fuming stage by supplying some reducing agent, such as coal or hydrocarbon, whereby a reducing gas containing, for example, CO, CO2 and metallic zinc is exhausted. Slag is then discharged from first fuming stage.
The invention also relates to an apparatus for treating zinc concentrates, the apparatus including a smelting furnace and a slag treating furnace arranged in connection therewith. A partition wall separates gas spaces of the smelting and slag treating furnaces from each other, and another partition wall separates the slag layers from each other.
Today, almost all zinc is produced by a hydrometallurgical process, i.e. by electrolysis, or by a pyrometallurgical process in accordance with the ISP, i.e. Imperial Smelting Process.
According to the electrolytic process, which is usually used for zinc concentrates with a high content of zinc, the zinc oxide is leached directly out of a calcined zinc concentrate. The solution of zinc is purified, and zinc is recovered from the purified solution by means of electrolysis, whereby metallic zinc is precipitated eletrolytically. For a good yield of zinc, iron must be leached out and then precipitated, usually as jarosite or goethite. Jarosite, however, poses a great waste problem, partly because it is produced in large amounts and must be stored, and partly because it may contain Zn and possibly Cd, both of which, at least in large amounts, are considered toxic. As a result, it has become more and more difficult to obtain permission to store any significant quantity of jarosite.
Various ways have been explored for changing jarosite to a form in which it could be stored without causing any harm to the environment. Different hydrometallurgical methods have been suggested. For example, leaching processes which produce iron oxide with fine particles have been recommended. However, it is difficult to find a practical use for iron oxide comprised of fine particles. Smelting processes for producing an end product suitable for storing have also been suggested. Drying and smelting of jarosite would, however, call for a considerable heat volume.
In accordance with the pyrometallurgical ISP process, it is also possible to recover zinc from ore concentrates having a relatively low zinc content. The process involves two stages. First, the zinc concentrate is treated in a sintering or calcinating stage in order to separate SO2 and to oxidize Zn to ZnO. In a second, separate stage, the product containing zinc oxide is smelted in a smelting furnace. Coke is used in the smelting process as both fuel and reducing agent. Entrained with the reduction gases, metallic zinc is discharged from the smelting furnace and is usually retained in a splash condenser. The zinc recovered from the ISP process, however, is not of as high a quality as the zinc recovered from the eletrolytic process.
In PCT application WO 87/03010, it has also been suggested to recover zinc by direct oxidizing and smelting of zinc concentrate to metallic zinc. According to that method, smelting, oxidizing and reducing are effected in a two-stage furnace, where a smelting furnace and a fuming furnace together form a unit. The furnace is divided into a smelting section with an oxidizing zone and a fuming section with a reducing zone by means of a vertical partition wall which divides the upper section of the furnace into two parts. The partition wall is so arranged in the furnace as to effect transport of slag between the various zones under the partition wall. Zinc concentrate and oxidizing gas are introduced into the smelting furnace or the oxidizing zone by means of lances in order to slag zinc to zinc oxide and to drive off SO.sub.2.
The slag containing zinc oxide is conducted, below the partition wall, from the oxidizing zone to the fuming furnace for reduction of the slag. A reducing agent is introduced into the reducing zone by means of lances, whereby zinc oxide is reduced to metallic zinc, which is discharged from the furnace with the exhaust gases.
U.S. Pat. No. 4,741,770 discloses a method of recovering zinc from zinc concentrate containing iron. This takes place in a furnace which is divided into a smelting section having an oxidizing zone, and a fuming section having a reducing zone. Zinc concentrate slags to produce zinc oxide in the oxidizing zone and is thereafter conducted to the reducing zone. In the reducing zone, ZnO is reduced to Zn, which is exhausted with the flue gases. The flue gases and the slag in the reducing zone are provided with heat energy by means of combusting part of the reducing gases remaining in the reducing zone directly above the slag surface. Iron containing slag is tapped off of the reducing zone.
According to the method disclosed in the above-mentioned US patent, part of the slag discharged from the fuming section may be conducted to an additional reduction stage, where part of the iron contained in the slag is reduced. The reduced iron is recirculated to the fuming section where it contributes to the reduction of ZnO to Zn. The slag from the additional reduction stage may then be led to the smelting furnace in order to facilitate the slagging of the ferrous zinc concentrate.
In spite of direct circulation of smelted slag from the smelting furnace to the fuming furnace, utilization of the iron contained in the ore concentrate for reduction of ZnO, and recirculation of slag from the additional iron reduction stage to the smelting furnace, the energy requirement of the processes described above is relatively high.
Zinc recovery from ore concentrates rich in zinc involves a problem with slagging because the zinc content of zinc slag cannot exceed 25%; otherwise its viscosity becomes too high. Addition of further slag components for maintaining the zinc concistency at a low level for obtaining a maximal amount of zinc oxide from the slag calls for more energy for heating and smelting the slag components.
An object of the present invention, therefore, is to provide a method for recovering zinc from zinc concentrate, in which method the need for reducing agents and fuel, such as coal or hydrocarbon, is considerably lower in comparison with the prior art methods described above.
Another object of the present invention is to provide a method of producing a slag which is easier to store.
A still further object of the present invention is to provide a simple and compact apparatus for recovering zinc from zinc concentrate.
An exemplary method of recovering zinc according to the invention is characterized in that part of the slag discharged in the first fuming stage is directly recirculated to the smelting and oxidizing stage in order to recover zinc oxide from the slag in the smelting and oxidizing stage. The slag formed in the smelting and oxidizing stage is preferably reduced in two successive fuming stages so that
slag from the smelting stage is reduced in the first fuming stage to a zinc content of 5 to 15%, preferably &gt;10%;
part of the discharged, reduced slag from the first fuming stage is recirculated to the smelting and oxidizing stage; and
another part of the discharged, reduced slag is further led to a second fuming stage, where the slag is reduced to a zinc content &lt;5%, and preferably &lt;2%.
The apparatus for separating zinc from zinc concentrate according to the invention is characterized in that
the slag treating furnace is divided into an inlet zone and an actual fuming zone by means of a partition wall extending downwardly from the ceiling of the furnace, the partition wall being so arranged that the upper gas spaces and the upper slag layers of both the inlet zone and the fuming zone are separated from each other, and that the inlet zone and the fuming zone are in communication mainly in the lowest part of the furnace;
an overflow is arranged in the partition wall between the smelting furnace and the inlet zone so as to facilitate slag flow from the smelting furnace, where the slag level is higher, to the inlet zone; and
a second overflow is arranged in the partition wall between the actual fuming zone and the smelting furnace so as to facilitate flowing of reduced slag to the smelting furnace from the fuming zone, where the slag level has been raised over the slag level in the smelting furnace, by injecting air, coal or hydrocarbon.
The zinc concentrate, i.e., zinc sulfide concentrate may be oxidized and smelted in a flash furnace, cyclone furnace or some other furnace suitable for smelting of ore concentrate. The smelted slag containing zinc oxide may be led directly to a continuous process in a slag treating furnace for reduction of zinc to metallic zinc, whereby the slag may be transported by means of a pouring ladle from the smelting furnace to the slag treating furnace. This process may also be effected in batches.
The problem arising from too high slag viscosity with high zinc content when zinc is recovered from a concentrate rich in zinc is avoided in this invention by recirculation of slag poor in zinc from the slag treating furnace to the smelting stage. The slag poor in zinc lowers the zinc consistency of the slag in the smelting stage, whereby the slag maintains its viscosity. Addition of ready-smelt hot slag does not require any additional energy for heating or smelting. Slag may be recirculated to such an extent that a so-called saturation limit of zinc is achieved. The saturation limit is determined by the viscosity of the slag. A majority of zinc concentrates have the saturation limit in about 25% zinc consistency. Slag from the slag treating furnace may be transported either continuously to the smelting furnace or by means of a pouring ladle.
Fuming i.e. reduction of Zn from the slag, may be effected in a conventional fuming furnace, electric furance, a furnace provided with lances or some other type of furnace suitable for a reduction process.
The zinc is discharged from the slag treating furnace with the exhaust gases in the form of volatile metallic zinc. The zinc may be oxidized to ZnO and be separated eletrolytically or be condensed directly from the exhaust gases in a splash condenser.
According to the invention, the need for a reducing agent and fuel for the zinc process is considerably decreased by means of effecting the zinc oxide reduction in two stages. In the first stage, the slag is reduced to a zinc consistency of 5 to 15 %, preferably to about 10%. Reduction to a zinc consistency of 5 to 10% is very fast and easy to perform, and the need for reducing agents is considerably lower with a high zinc content than with a zinc content below 5%.
Part of the slag from the first reduction stage is recirculated to the smelting furnace and only, for example, 25 to 50% of the slag is led further to a second reduction stage, to be reduced there to a zinc content &lt;2%. The amount of slag in the second reduction stage is considerably lower than in the first reduction stage. The amount of slag in the second reduction stage is a function of the Fe or SiO.sub.2 content of the ore concentrate. Since only a little amount of slag is reduced to a low zinc content, the need for reducing agents is much lower than if the whole slag amount from the first reduction stage were reduced to a low zinc content. An inert silicate slag is recovered from the second reduction stage, i.e. a slag which does not easily react and which is therefore easy to both store and utilize for various purposes.
In accordance with a preferred embodiment of the invention, transport with pouring ladles can be avoided by combining the smelting furnace and the slag treating furnace into one unit. The slag treating furnace is so constructed as to provide continuous recirculation of slag from the fuming zone into the smelting furnace and transport of slag from the smelting furnace to the inlet zone of the slag treating furnace.
It will therefore be appreciated that, according to the invention, a process consistent with enviromental requirements has been developed, which also provides advantages in terms of energy efficiency. If the invention is applied to a process with electrolytic zinc recovery, the process will be simpler because the acid leaching stage and the iron precipitation stage are no longer necessary since the iron remains in the slag. A further advantage of the novel method according to the invention is that the slag received is inert and easy to store. The method of recovering zinc according to the invention also consumes less coal or hydrocarbon than the ISP process. In addition, concentrates rich in zinc ore may be utilized without any need for adding large amounts of slag-forming components.
Additional objects and advantages of the invention will become apparent from the detailed description which follows.